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GEOMECHANICAL EVALUATION OF A MECHANICAL UNDERCUTTING SYSTEM IN

BLOCK CAVING

 by

Fernando Villegas

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A thesis submitted to the Faculty and the Board of Trustees of the Colorado School of

Mines in partial fulfillment of the requirements for the degree of Doctor of Philosophy (Mining

and Earth Systems Engineering).

Golden, Colorado

Date: ______________________

Signed: ______________________________Fernando Villegas

Signed: ______________________________Dr. Mark Kuchta

Thesis Advisor

Golden, Colorado

Date: ________________________

Signed: ______________________________Dr. Pricilla P. Nelson

Professor and HeadDepartment of Mining Engineering

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ABSTRACT

 Nowadays, Block Caving is the underground mining method with the highest productivity in

the world. Many of the components of block caving have changed in the last few years, as new

techniques and equipment have become available, increasing the efficiency of the method. Such

components include automatic LHD (load, haul and dump) trucks and trains, electronic

detonators and mechanical equipment for the development of drifts and shafts.

 Nevertheless, the first link of the chain of value in Block Caving is the undercutting process,

which continues, in essence, being performed in the same way it was when the method was

invented, utilizing drill and blast. This thesis proposes to use a mechanical system to perform the

undercutting, similar to the longwall method used in coal mines. In order to evaluate the

technical feasibility of this option, a geomechanical evaluation about the feasibility of theimplementation of this alternative system in block caving was developed.

Starting with the geomechanical study, the first step was exploring the possible

geomechanical hazards that could arise if this alternative method is implemented in block caving,

taking into consideration that the undercutting work should be performed from inside of caving

volume. An undercutting mine design was proposed and interest zones were defined, also a risk

analysis was carried out. The result of the risk analysis was the identification of three possible

hazards, losses of confinement in the cutting front, wedges in the roof of the work area, and

collapse of the pillar between panels.

In order to assess the geomechanical hazards, numerical models and kinematic analyses were

conducted. For these, numerical models were built in FLAC3D  to study the geomechanical

 behavior of the cutting front, work area, and the pillar between panels. The models further

studied the relation between each of these elements and the undercutting advance and caving

 propagation. A mine design and sequence has been proposed for these analyses. The mine design

was made based on the macro block method and the sequence of undercutting advance was made

 based on the longwall method.

The undercutting front advance was modeled in two stages, pre-caving and continuous

caving, in order to take into consideration the undercut advance and the caving progress. The

results of the analyses showed that the stress quickly raised with the increase of the undercut area

until it reached the hydraulic radius of caving. In order to estimate the initiation and depth of

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fracturing in the cutting front and to evaluate the risk of losses of confinement in the center of the

cutting front, two methodologies were used: UCSlab and critical strain εc. These methodologies

showed that on the cutting front fracturing started when the undercut area reached 20 to 30% of

the panel area. Thus, the conditions where the equipment will work change constantly with the

advance of the undercutting front. Therefore, the machine should cut and remove the rock in the

cutting front. In addition, three different cutting front widths were analyzed: 120, 150, and 180

meters. For this range of widths, the difference in stress behavior is low. On the other hand,

estimates of vertical displacements in the work area increase with the advance of the

undercutting front, but near the cutting front the vertical displacements will be low; therefore,

stability in the work area could probably be controlled structurally by wedges formed in the roof.

 Numerical models show that during continuous caving, the length of the beam does not have

greater influence on variations in the stress condition. Regarding the pillar between panels, it

should accomplish the function of giving continuity to the method and helping it to provide

stability of the work area. The pillar design is robust, with a width-to-height ratio of five;

therefore, it has a high strength. The numerical models confirm this approach by showing that

although there is damage in the sides of the pillar, the damage never reaches the center of the

 pillar.

Estimation of the possible undercutting advance rate using a mechanical undercutting system

in block caving shows that it is similar to or higher than the traditional drill and blast rate.The research developed in this thesis presents an opportunity to improve the quality of

undercutting, ensuring the cut of the block and eliminating the risk of a remnant pillar that would

affect the production level. In addition, it delivers the option of a possible increase in the

 performance of the undercutting, and an exploration of a possible automation of the process, both

of which would contribute to the modernization of block caving exploitation. The main

conclusion of this work is that from a geomechanical point of view, it is feasible to use a

mechanical undercutting system in block caving.

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TABLE OF CONTENTS

ABSTRACT ........................................................................................................................ iii 

LIST OF FIGURES ..................................................................................................................... ix 

LIST OF TABLES .......................................................................................................................xv 

ACKNOWLEDGEMENTS ...................................................................................................... xvi 

INTRODUCTION ..........................................................................................1 CHAPTER 1

1.1 

Background ......................................................................................................1 

1.2  Justification for the research conducted ...........................................................2 

1.3 

Objective of the research undertaken ...............................................................2 

1.4  Research conducted and methodology .............................................................3 

1.5  Scope of the research ........................................................................................4 

1.6 

Original contribution ........................................................................................4 

1.7  Thesis organization ...........................................................................................5 

LITERATURE REVIEW ..............................................................................7 CHAPTER 2

2.1  Block caving method ........................................................................................7 

2.1.1  Conceptual model of caving .........................................................7 

2.2  Undercut methods .............................................................................................9 

2.2.1  Post undercut .................................................................................9 

2.2.2  Pre undercut ................................................................................10 

2.2.3 

Advance undercut .......................................................................11 

2.2.4 

“Crinkle cut” layout and blasting design in advanceundercutting ................................................................................12

2.2.5  Cavability ....................................................................................13 

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2.2.6  Undercutting rate ........................................................................14 

2.2.7  Geometry of undercutting front ..................................................15 

2.2.8  Undercutting method in high stress ............................................16 

2.2.9 

Generation of dynamic stress ......................................................18 

2.2.10  Seismicity associated to undercutting process ............................20 

2.2.11   Numerical models of undercutting in block caving ....................20 

2.3  Longwall mining method ...............................................................................21 

2.3.1  Mine design .................................................................................22 

2.3.2 

Production system .......................................................................23 

2.3.3   Numerical model of longwall .....................................................25 

2.4 

Chuquicamata underground project ...............................................................27 

2.4.1  Chuquicamata mine ....................................................................27 

2.4.2  Geology .......................................................................................29 

2.4.3  Mining method ............................................................................30 

UNDERCUTTING WITH A MECHANICAL SYSTEM ........................33 

CHAPTER 3

3.1 

Block caving mine designs with mechanical undercutting ............................33 

3.1.1  Undercutting mine design ...........................................................34 

3.1.2 

Relation between undercut design and production leveldesign ..........................................................................................38

3.2 

Undercutting sequence ...................................................................................38 

3.3  Performance of mechanical undercutting system. ..........................................39  

3.3.1  Productivity in advance undercut with drilling and blasting ........................................................................................39

3.3.2 

Advance undercut with a mechanical undercutting system ........40 

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3.4  Possible differences between the use of a mechanical undercuttingsystem in block caving and longwall ............................................................44

GEOMECHANICAL HAZARDS IN THE USE OF A MECHANICALCHAPTER 4

UNDERCUTTING SYSTEM ......................................................................45

4.1  Methodology used for risk analysis ................................................................45 

4.2 

Description of hazard .....................................................................................46 

4.2.1  Cutting front ................................................................................46 

4.2.2 

Work area ....................................................................................48 

4.2.3  Drifts access to panel ..................................................................50 

4.2.4 

Pillars of separation between panels ...........................................51 

4.2.5  Caving volume ............................................................................52 

4.3  Evaluation of risks ..........................................................................................58 

NUMERICAL MODELS .............................................................................61 CHAPTER 5

5.1  Framework of numerical models ....................................................................61 

5.1.1 

 Numerical models objectives ......................................................61 

5.1.2   Numerical model software ..........................................................62 

5.1.3  Caving condition .........................................................................62 

5.2  Modeling with FLAC3D ..................................................................................63 

5.3  Modeling construction ....................................................................................66 

5.3.1   Numerical model geometry .........................................................67 

5.3.2 

Boundary conditions ...................................................................72 

5.3.3 

Stress condition ...........................................................................72 

5.3.4  Constitutive model ......................................................................74 

5.3.5  Rock mass properties ..................................................................74 

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5.3.6  Meshing and element size ...........................................................76 

5.4  Excavation sequence ......................................................................................77 

5.4.1  One Panel ....................................................................................77 

5.4.2 

Two Panels ..................................................................................79 

5.5  Criteria to estimate the initiation and depth of fracturing in the cuttingfront ...............................................................................................................82

5.5.1 

Estimation of initiation and depth of fracturing in thecutting front based on the uniaxial compression strengthof intact rock (UCSlab) ................................................................82

5.5.2 

Estimation of depth fracturing in base of critical strain ..............84 

GEOMECHANICAL ANALYSES.............................................................86  CHAPTER 6

6.1 

Framework of geomechanical analyses ..........................................................86 

6.2  Geomechanical behavior of cutting front and work areas during pre-caving ............................................................................................................86

6.3 

Geomechanical behavior of cutting front and work areas incontinuous caving .........................................................................................91

6.4  Effect of the width of the panel in the cutting front .......................................94 

6.5  Stability of the pillar between panels .............................................................95 

6.6  Kinematic analyses of the work area ............................................................101 

6.7   Numerical model validation .........................................................................106 

CONCLUSION AND FUTURE WORK  ..................................................110 CHAPTER 7

7.1 

Conclusions ..................................................................................................110 

7.2 

Future work ..................................................................................................112 

REFERENCES CITED .............................................................................................................113 

APPENDIX A  CODES FLAC3D OF NUMERICAL MODELS ......................................121 

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LIST OF FIGURES

Figure 2.1 Typical scheme of block caving with LHD (Arce, 2002) ..............................................7

Figure 2.2 Conceptual model of caving (Brown, 2007) ..................................................................8

Figure 2.3 Post-undercut (Brown, 2007) .......................................................................................10

Figure 2.4 Pre-undercut (Brown, 2007) .........................................................................................11

Figure 2.5 Advanced undercut (Brannon, 2011) ...........................................................................12

Figure 2.6 Illustration of crinkle cut in Palabora mine (Brown, 2007)..........................................14

Figure 2.7 Crinkle cut undercut potential problem areas (Butcher, 2000) ....................................14

Figure 2.8 Relative frequency of different undercut rates in caving mines (Flores, 2002) ...........15

Figure 2.9 Confinement and traction in the undercutting front (modified from Laubscher,2000) ............................................................................................................................16

Figure 2.10 Geometries of the undercutting front (Butcher, 1999) ...............................................16

Figure 2.11 Variation of in-situ stress with depth (Karzulovic, 2004) ..........................................17

Figure 2.12 Stress conditions in the roof of production drift by effect of the undercuttingfront advance (Trueman, 2002) ................................................................................17

Figure 2.13 Example of advanced undercutting front with small pillars (Workshop

undercutting, 2008) ...................................................................................................18

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Figure 2.14 Remnant pillar in undercut level (Karzulovic, 2005) .................................................19

Figure 2.15 Damage caused by collapse in production level (Karzulovic, 2003) .........................19

Figure 2.16 Cumulative undercut area blasted (green dots) compared to the cumulativenumber of significant events (red line) showing a very strong correlation(Potvin, 2010) ...........................................................................................................20

Figure 2.17 Classic schematic of longwall, in a coal mine (Energy InformationAdministration, 1994) ...............................................................................................22

Figure 2.18 A typical longwall coal mine layout showing the panels, entries, and chain pillars, (Source: Ch13.8, SME Handbook, 2011). ....................................................24

Figure 2.19 Overview of longwall system, main equipment (University of Wollongongwebsite, 2014) ...........................................................................................................25

Figure 2.20 Three dimensional model (Wei, 2011) .......................................................................26

Figure 2.21 Chuquicamata Underground Mine, location (Fuentes, 2009) ....................................28

Figure 2.22 Chuquicamata mine (Fuentes, 2009) ..........................................................................28

Figure 2.23 Macro blocks and geology, first level, Chuquicamata Underground Project(Villegas, 2008) ........................................................................................................29

Figure 2.24 Ore deposit characterization (Fuentes, 2009) .............................................................30

Figure 2.25 Position of production levels in Chuquicamata Underground Project(Fuentes, 2009) .........................................................................................................31

Figure 2.26 Macro block method (Fuentes, 2009) .........................................................................31

Figure 2.27 Macro block configuration in first and second levels of ChuquicamataUnderground Project (Allende, 2009) ......................................................................32

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Figure 3.1 Differences in location of the work area in classical undercutting andmechanical undercutting system ..................................................................................33

Figure 3.2 Scheme of mechanical undercutting system in block caving (modified fromBrown, 2007) ...............................................................................................................34

Figure 3.3 Layout undercut level for a mechanical undercutting system ......................................36

Figure 3.4 Section layout undercut and production level in a mechanical undercuttingsystem ..........................................................................................................................36

Figure 3.5 Plan view of undercut mine design for mechanical undercutting system in block caving .................................................................................................................37

Figure 3.6 Illustration of mechanical undercutting system ............................................................37

Figure 3.7 Mine design for mechanical undercutting system in block caving ..............................38

Figure 3.8 Concept of drum miner (Frenzel, 2011) .......................................................................43

Figure 5.1 Laubscher’s caving chart (Brown, 2007) .....................................................................63

Figure 5.2 The basic calculation cycle of FLAC 3D (Itasca, 2012) ..............................................64

Figure 5.3 Numerical model process (Itasca, 2012) ......................................................................68

Figure 5.4 Process map of numerical model in block caving undercutting (Modify fromItasca 2012). .................................................................................................................69

Figure 5.5 Plan view of numerical model mesh for one panel ......................................................70

Figure 5.6 Configuration of model for one-panel FLAC3D ..........................................................70

Figure 5.7 Geometry of numerical models two panels ..................................................................71

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Figure 5.8 Configuration of model for two-panel FLAC3D ..........................................................71

Figure 5.9 Numerical model mesh for two panels .........................................................................72

Figure 5.10 Border conditions in the numerical models ................................................................73

Figure 5.11 Theoretical model of strain-softening (Cai, 2007) .....................................................75

Figure 5.12 Increase in the mesh size. ...........................................................................................77

Figure 5.13 Model of undercutting sequence for one panel ..........................................................78

Figure 5.14 Isosurface of vertical displacement to find the steps of undercut for thenumerical model. ........................................................................................................78

Figure 5.15 Sketch sequences of numerical models for continuous caving ..................................79

Figure 5.16 Sequence of modelling of the pillar between panels ..................................................80

Figure 5.17 Plot of laboratory tests in PEK (Board, 2006) ............................................................82

Figure 5.18 Empirical thresholds for in-situ damage in moderately jointed hard rock mass(Diederichs, 2003) ......................................................................................................83

Figure 5.19 Concept of initiation of fracture and slabbing (Stacey, 1981) ....................................85

Figure 6.1 Scheme of analysis of stress in the cutting front ..........................................................87

Figure 6.2 Behaviour of vertical stress with advance of the undercutting front ............................88

Figure 6.3 Behaviour of normal stress with the advance of the undercutting front .......................88

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Figure 6.4 Behaviour of horizontal stress with the advance of the undercutting front ..................89

Figure 6.5 Increase of displacements with the advance of the undercutting front ........................90

Figure 6.6 Variation of depth of fracture in the cutting front with the advance of theundercutting front.........................................................................................................91

Figure 6.7 Prediction of fracturing in the cutting front with 120 meters of undercuttingfront advance with UCSlab and critical strain criteria .................................................92

Figure 6.8 Distributions of tress in the cutting front of different undercutting advanceoperations in continuous caving with a 40 meters beam .............................................93

Figure 6.9 Distributions of vertical stress in the cutting front for different lengths of beams with an undercutting advance of 180 meters ....................................................93

Figure 6.10 High vertical displacement in the work area with a 20 meters beam .........................94

Figure 6.11 Vertical displacements in the cutting work area with different sizes of beams .........95

Figure 6.12 Variation of principal stress with the undercutting front advance and changeof the cutting front condition ....................................................................................96

Figure 6.13 Variation of stress in the undercutting front for different widths of theundercutting front .......................................................................................................96

Figure 6.14 Sequence for modelling of Panel B and pillar ............................................................97

Figure 6.15 Mining steps in the life of the pillar between panels ..................................................98

Figure 6.16 Monitoring of vertical stress in the middle of the pillar .............................................98

Figure 6.17 Profile of vertical stress in the pillar to 100 meters of undercutting advance ..........101

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Figure 6.18 Profile of vertical stress in the edge of the pillar to the position of theundercutting front .....................................................................................................102

Figure 6.19 Fracturing zones in the pillar using UCSlab criterion ..............................................103

Figure 6.20 Estimated fragmentation of macro blocks for Chuquicamata UndergroundProject, using BCF (Board, 2006) ............................................................................104

Figure 6.21 Three-dimensional view of maximum wedge in work area, and oneillustrative example (Brown, 2007) ..........................................................................104

Figure 6.22 Plan of view of maximum wedge and hydraulic roof support .................................105

Figure 6.23 Possible wedges in the border of the beam and the caving volume .........................105

Figure 6.24 Brittle failure in longwall face (Qing- Sheng, 2013) ................................................106

Figure 6.25 Estimation depth fracture in the cutting face from Thesis work ..............................107

Figure 6.26 Stress redistribution around a coal mining face (Qing-Sheng, 2013) ......................107

Figure 6.27 Result of numerical models stress distribution (Figure 6.2).....................................107

Figure 6.28 Collapse of the hanging wall of the working for a sub-critical support pressure (Gonzalez, 2008) ........................................................................................108

Figure 6.29 Plasticity zones for 20 and 40 meters undercutting advances, showing theinstability zones in the roof ......................................................................................109

Figure 6.30 Numerical model fracturing initiation and exploration tunnel condition. ................109

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LIST OF TABLES

Table 3.1 Productivity estimation for an advance undercut with drill and blast. ..................... 40 

Table 3.2 Muck to be removed in advance undercut. ............................................................... 40 

Table 3.3  Parameters for estimating the productivity of a mechanical system withdifferent widths of the undercutting front. ................................................................42 

Table 3.4  Estimate of productivities for different widths of undercut fronts. .......................... 43 

Table 4.1 

Matrix of evaluation of probability of occurrence (likelihood) ................................ 45 

Table 4.2  Matrix of evaluation of impact (consequence) ......................................................... 45 

Table 4.3 

Matrix to identify the significant risks of the project ............................................... 46 

Table 4.4 Identification of geomechanical hazards during the undercutting processusing a mechanical undercutting system ...................................................................54

Table 4.5 Matrix of risk assessment ......................................................................................... 58 

Table 5.1 Stress condition of macro block N-42 (see Chapter 2) level 1841 ........................... 73 

Table 5.2 Properties of intact rock (Laboratory tests, Russo 2010) ......................................... 75 

Table 5.3 Properties of rock mass (Russo, 2010) ..................................................................... 76 

Table 5.4  Consideration of numerical models .......................................................................... 81 

Table 5.5 Detail of numerical models built .............................................................................. 81 

Table 6.1  Structural sets in macro block N 42 (Zaro, 2009). ................................................. 103 

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ACKNOWLEDGEMENTS

I would first like to express my gratitude to my first and second advisers in the mining

department of Colorado School of Mines, Dr. Christian Frenzel and Dr. Mark Kuchta. To

committee member Dr. Gideon Chitombo, who accepted a place on this committee and

supported me from Australia, I offer my thanks for his friendship and advice. I also thank Dr.

Gubb, Dr. Nakagawa, and Dr. Cohen for their help on my research.

I would like to acknowledge God and my family for their support, motivation, prayers, and

love: my wife Fresia and my children Fernando, Francisca, and Pía. Also, I would like to thank

my mother and brother for their constant concern and support.

I am very grateful to my company, Codelco-Chile, which was my sponsor during this time — 

especially the person that encouraged me to study and supported me throughout, Mr. SergioFuentes. Thanks also to my colleagues Samuel Peñaloza and Andrea Russo.

I thank Itasca Consulting group in Minnesota for helping me to use FLAC 3D, with special

thanks to Dr. Loren Lorig, Thierry Lavoie, Math Pierce, David DeGagne, and my friend Mark

Board and his wife Cathy.

I am very grateful to my friends Matias, Beatriz, Andres, and Adalberto. Thank you for your

friendship. And thanks to my English teacher, Joyce, for her help and sympathy.

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CHAPTER 1

INTRODUCTION

1.1  Background

The largest mining companies in the world have begun the construction phase of their biggest

 projects in block caving mining history. In the coming decades we will see that underground

 production rates increase more dramatically than ever before (Brown, 2007 b, Chitombo, 2010).

Block caving is an efficient method of underground mining with a high production rate; it is the

method currently used in the biggest underground mines around the world. It has evolved

through time from a manual to a mechanized production operation, adapting to different rock

qualities and stress conditions. During the application of block caving, the first activity in the

 process to generate the caving is the undercut, which uses drilling and blasting to cut the base ofthe blocks (Hustrulid, 2001).

Modern block caving operations such as Palabora, DOZ, and El Teniente have incorporated

changes in the sequences of mining construction. Most of these changes have been related to the

undercutting advance, because this process causes an important increase in the abutment stress,

which in turn damages the infrastructure near the undercutting front (Chitombo, 2010). In order

to mitigate this effect, new undercutting variants have been incorporated which are pre-

undercutting and advanced undercutting (Bartlett, 2000). These changes have led to an increase

in the available area for the production and a decrease in the repair rates for the production levels

(Araneda, 2007). However, along with these positives come negative effects in the undercut

level, because these variants need around twice as much development in the undercut level,

compared with the post undercut method, to compensate for the swelling factor and achieve an

adequate undercutting blasting. This causes a decrease in the width of the pillars in the undercut

level, which leads to reduction in strength of the pillars, causing collapses in undercut and

 production levels (Villegas, 2008).

 Nowadays, considering the deepening of mining (and high stresses due to the height of the

rock column), the pre-undercutting and advanced undercutting variants would be the most

adequate for a high stress condition. Nevertheless, the stability problem in the undercut level

(specifically the pillar collapses due to the thinness of the rock pillars) has not yet been solved,

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and it is necessary to make mitigation efforts. Such efforts may include pre-charge of explosives

or use of extra support in the undercut level (Bartlett, 2000). 

Due to a shortage of new mineral deposits close to the surface, the trend of the mining

industry is toward underground exploitations, and block caving is the method most used and

which provides the highest production (Brown, 2007).

One of the main concerns during mining by block caving is to avoid leaving areas with a

deficient undercut, which could provoke intense damage in the levels below and generate

collapses that result in loss of productive areas. In contrast, the longwall caving method that is

used in underground coal mining has overcome this problem. The longwall method uses a

machine to cut the rock, a series of hydraulic shields to maintain ground stability in the work

area, advancing and cutting the base of the rock column. The method provides absolute certainty

that the cut of the panel is total, without the possibility of a remnant pillar (Peng, 2006).

1.2  Justification for the research conducted

Codelco is in the first phase of construction of the two biggest underground projects in the

world: the Chuquicamata Underground Project and the New Mine Level, name given to the new

 production sector in El Teniente mine. They will be exploited by using the block caving method,

with a production of 140,000 ton per day. The topic developed in this thesis is part of the aim of

modernization of future operations through changes in the process. In this case, changes to the

undercutting process are examined in order to ensure the cut quality and research of possible

management opportunities from changing a discontinuous undercutting process such as drill and

 blast to a continuous process such as mechanical undercutting, eliminating or mitigating the

negative geomechanical effects. Based on the aforesaid, this thesis proposes the possibility to use

longwall method as a mechanical undercutting system. This is based on the experience of the

longwall method in coal mining

1.3 

Objective of the research undertaken

The main objective of this thesis is to evaluate, based on a geomechanical point of view, the

technical feasibility of using a mechanical excavation system for undercutting in underground

 block caving.

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This thesis presents a geomechanical evaluation of a proposed alternative method for the

classical undercutting system. Advantages of the proposed undercutting system include ensuring

the undercutting and decreasing the quantity of development in the undercut level. This in turn

would improve excavation stability and reduce long-term costs. The use of a mechanical system

would also offer the ability to increase the undercutting rate and offer the possibility to automate

the system.

In the first part of this work the possible geomechanical aspects of the use of a mechanical

undercutting system in block caving are identified and evaluated. The objective of this new

approach may be applied in future block caving operations. During this first part, it was

necessary to explore all possible stability problems in different stages of the caving, and identify

the places more susceptible to damage. Every condition is explained and illustrated for a better

understanding. With this information, a risk evaluation was made in order to identify the mostcritical problems for the geomechanical analyses. This work should also be useful for possible

industrial testing.

The geomechanical evaluations of the main risks were made through analytical and

numerical models, to identify parameters and conditions that control the stability problems and

the alternative solutions. The numerical tool used for the geomechanical analyses was FLAC3D.

To develop this thesis, the data of the Chuquicamata Underground Project was used.

The numerical models were used in order to estimate the stress behavior in the undercutting

front when the undercut front advance and the caving propagation grow, all within a scheme of

advance undercutting similar to the longwall method.

A resulting consideration is the evaluation of the technical feasibility of controlling possible

stability problems in the area of the operation of the cutting system, and in the pillar between

undercut panels, in order to permit a safe operation for equipment and people.

Another objective is to present a rough estimation of possible productivity of a mechanical

undercutting system compared to the classical drill and blast.

1.4  Research conducted and methodology

The research was conducted based on a geomechanical point of view, an evaluation of the

 possible application of a mechanical undercutting system in block caving. The bibliographic

review was focused on geomechanical issues occurred during the undercutting process with

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different undercutting methods, this due to that there are not experiences in applying mechanical

undercutting in block caving. Because of this, it was necessary to propose a type of undercutting

method with a mine design for mechanical undercutting. Based on the review and the Codelco

Chile experience in block caving operations, an identification and evaluation of the possible risks

in the implementation of a mechanical undercutting system were undertaken to recognize and

 prevent the most relevant hazards. Mining design and the rock mass properties used are derived

from the Chuquicamata Underground Project.

With the information previously mentioned, numerical models were completed by using

different caving stages and different widths of the undercutting front. The results of the

numerical models showed the stress and deformation, and the analyses resulted in obtaining a

 behavior hypothesis, which has made possible an estimation of the potential performance of the

system’s productivity and the exercise of a comparison with the drill and blast classical system.

1.5  Scope of the research

This work evaluates the technical feasibility of incorporating the mechanical undercutting in

the block caving process, from a geomechanical approach, without considering the mechanical

design of the equipment. Currently, the equipment — such as the shearer and hydraulic support — 

are developed for soft rock, but the technology associated with cutting tools and roof support

capacities have grown significantly in recent years. This study is offered as a starting point

toward the adaptation and development of this technology in block caving. Furthermore, the

mechanical design of the equipment is beyond the scope of this work.

The information used in the analyses and numerical models is from the Chuquicamata

Underground Project. This project’s  characteristics, in terms of rock properties and stress

conditions, correspond to average block caving operations elsewhere in the world. Also, this

 project sponsored the author of this thesis during the time of this research.

1.6 

Original contribution

 Nowadays, there is little research about undercutting system in block caving, and the latest

innovations associated with the undercutting, such as pre and advance undercut, have increased

the complexity of the system. In addition, the use of drill and blast system to initiate the caving

has never been questioned.

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This thesis is the first work to consider and evaluate, from a geomechanical point of view, the

application of a mechanical undercutting system in block caving.

In this work, there is a mine design and an undercutting sequence which would allow the

 possible implementation of a project for industrial testing. In addition, this thesis presents an idea

about of the stress behavior with the undercut, which could be used by mining equipment

manufacturers to design or adjust the technology necessary for this application in block caving.

Another original contribution is a description of the possible geomechanical risk in the

eventual uses of a mechanical system in block caving. This alternative, use a mechanical system

to undercutting, opens the door for automation of the undercut system in block caving.

This work is an innovation work, never before in the metallic mining industry has been

evaluate the option of use a mechanical system to do the undercutting in block caving, therefore

an objective of this thesis will be an original contribution to the mining engineering.

1.7  Thesis organization

This thesis is divided into seven chapters.

Chapter 1 is an introduction that explains the motivation of the study and the problem

statement, current geomechanical challenges in block caving, the objectives of the study and the

scope of the thesis and how it is organized. Also, original contributions of this study are

discussed and suggestions for further research are provided.

Chapter 2 provides a literature review, which present the different block caving undercut

alternatives, the longwall method in coal mines and an overview of the study sector which was

used as the basis for the studies and analyses developed in this work.

Chapter 3 in this chapter there is a description of mining method proposed for the analyses

developed in this study, the mine design and performance for both methodologies are discussed,

drill and blast method and mechanical undercutting, from the perspective of the implementation,

and also gives a practical breakdown of the differences between a mechanical system and the

classical system of drilling and blasting.

Chapter 4 identifies and evaluates the possible geomechanical aspects in the use of a

mechanical undercutting system in block caving, and reveals the places most susceptible to

suffer damage. Based on this a risk analysis was done.

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Chapter 5 describes the organization of numerical models used for the studies and analyses of

the mechanical undercutting system, and also the criteria for the evaluations used.

Chapter 6 discusses and analyzes the results of the numerical models for the different steps of

the caving process in order to understand the behavior of the rock mass with regard to the results

of the undercutting process.

Chapter 7 provides a summary of the work undertaken and the conclusions drawn from this

study.

.

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CHAPTER 2

LITERATURE REVIEW

2.1  Block caving method

Block caving is a method of mining normally used in massive orebodies, in which a block of

rock mass is defined, this block is fully undercut to initiate the caving. The undercutting process

is performed through drill and blasting, the broken material is drawn to create a void to initiate

the caving. The broken material is removed through the production level from the drawbells.

Block caving is a low cost production method with high productivity, but requires

considerable investment in mine preparation. Today, there are large mining operations that use

mechanized block caving and some other operations have trials for automated mode Figure 2.1. 

Figure 2.1 Typical scheme of block caving with LHD (Arce, 2002)

2.1.1 Conceptual model of caving

Duplancic and Brady developed a conceptual caving model (Brown, 2007) Figure 2.2. The

model includes four zones each has different rock mass properties due to their different degrees

of fracturing. As caving progresses, the zones change sequentially in a process of rock mass

degradation associated with the propagation of the caving. The zones are defined as follows.

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o   Pseudo-continuous domain: The virgin rock mass around the caving, without the caving

effect. The behavior of the rock mass in general is elastic.

o  Seismogenic zone: In this zone, the geological discontinuities start to open and new

fractures are created. This generates seismic activity associated with damage to the rock

mass.

o   Zone of loosening : The rock mass in this zone has lost its cohesive strength. The rock

mass is intensely fractured, and only has residual strength.

o  Caving zone: This is fractured material that is to be removed through the production

draw. It does not have cohesion properties.

o   Air gap: The air gap is a zone that does not contain any material, which allows the caving

to progress.

Figure 2.2 Conceptual model of caving (Brown, 2007)

Based on this conceptual model of caving, the behavior of stress and strain in the rock around

the undercutting area is affected by the caving process. Therefore, the mine design is affected by

the caving geometry (which changes with the caving propagation and the undercutting advance).

Also, the size and the properties of the different zones defined in Figure 2.2 change, through a

degradation process caused by continuous redistribution of stress. To estimate the behavior of the

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stress/strain for any alternative or sequence of block caving using numerical models, it is

necessary to consider the conceptual model of caving.

2.2  Undercut methods

Caving is initiated with the undercutting process in which the base of the block or panel is

cut until the hydraulic radius of caving is reached (hydraulic radius is calculated dividing the

area by the perimeter of undercut zone). The material inside the caving zone is removed through

the drawpoints on the production level and the rock above this area collapses and the breaking

 process progresses until reaching the surface.

The undercutting activity is critical for the success of block caving (Laubscher, 2000), so it is

necessary to have a plan and a design for the undercut. If the cut is of low quality, it is possible

to lose production area, due to the damage or collapse of the production level, so it may benecessary to re-install ground support or rehabilitate some zones, which delays the production

schedule.

The traditional undercut system is post-undercutting. Recently developed undercutting

systems are the pre-undercut, advanced undercut, and macro block methods.

2.2.1 Post undercut

Post-undercutting is the classical undercut method, (Figure 2.3), all levels (undercut,

 production ventilation, transference, hammer, transport, and other) are developed before the

undercut process begins (drilling and blasting). Therefore, in this methodology it is necessary to

make investments in advance in order to build the necessary infrastructure. The advantage is that

when the undercut is done the sector is incorporated immediately into the production area. In this

method the pillars in the undercut level have more strength than that achieved with the pre-

undercut and advance undercut methods. The drifts in the undercut level have the same distance

 between them as the drifts in the production level, which gives greater pillar strength in the

undercut level. In addition, the undercut level is used only for undercut drilling and blasting, and

it is not necessary to handle muck on this level.

With post- undercuting, between one and three “drawbell lines” are blasted before the

undercutting front passes above the draw points on the production level, to absorb part of the

rock volume from the undercut blasting. Another advantage of post-undercutting is the relative

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 blasting. This has an important effect on the stability of the undercut level, which requires more

support than other methods such as pre-undercut and advance undercut. However, from an

economic point of view, this alternative has the advantage that investment in production level

infrastructure is made shortly before the area is incorporated into production.

Figure 2.4 Pre-undercut (Brown, 2007)

2.2.3 Advance undercut

In the advance undercut method, the undercut front advances ahead the blasting of drawbells

and part of the development and construction of the production level is done before the

undercutting front passes, Figure 2.5. This method has the advantage that it decreases the time

necessary to complete the development and construction of the production level. Also, like the

 pre-undercut method, it requires less support on the production level. Nonetheless, the

infrastructure developed before the undercut level needs strong support, and after the undercutfront has passed it is sometimes necessary to perform some reparation. This method increases the

effective life of the drawpoint, compared with post-undercutting. The disadvantage of this

method is the complexity of coordinating development, construction, and the advance of the

undercut front simultaneously. As with pre-undercutting, the advance undercut method requires

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twice the amount of development in the undercut level, to absorb the rock from the undercut

 blasting.

Figure 2.5 Advanced undercut (Brannon, 2011)

2.2.4 

“Crinkle cut” layout and blasting design in advance undercutting  

The narrow inclined undercut, or crinkle cut, corresponds to the advance undercutting most

used in block caving (Workshop undercutting 2008). It considers two concepts: the stability of

the undercut level and production level, and the clearance of muck generated from the blast

(clearance of muck creates the void necessary to contain the material for the next undercutting

 blast). To improve the stability of the production level, the height of the apex of the pillar

 between the drawbells is increased, and the drawbell is formed by the inclined blasting in the

undercut level (Butcher, 2000). In addition, the inclination helps to remove the muck because thematerial flows by gravity due to its own weight. Figure 2.6 illustrates the inclined blasting and is

 possible to see the drawbell, drilling design, blasting design and also the caving and production

levels all in 3D.

This design presents some problems inherent to the method (Casten, 2002). For example, it

creates the perfect condition to generate a coarse fragmentation in the caving initiation, due to

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the effect of the chevron geometry in the undercut roof. Butcher (2000) pointed out several other

 potential problem areas that must be taken into consideration when this method in used (Figure

2.7). These include: the possibility of leaving a remnant pillar in the apex of the crown pillar due

to poor drilling and blasting, muckpile stacking that affects the next blasting, and the requirement

of re-drilling in some blasting holes. Another difficult part of this method is the need for strict

control of the process: the crinkle cut requires strong discipline in the operation (Bartlett, 2000).

Moreover, the geometry of the undercutting advance (plan view) is irregular, with high stress

concentrations in the sharp parts, and the risk of losses of drills or abandonment of areas. Some

solutions for this problem have been the use of pre-charges of explosives and remote control

equipment to clear the area, in order to avoid the presence of people (Casten, 2002).

The undercutting method is highly demanding in the quantity of meters of the drift

development.In this method, the criterion is to open the drawbells behind the undercutting front, in a

distress zone with a distance that normally depends on the capacity to complete the infrastructure

of the production level (Araneda, 2007) (e.g., draw points construction, drilling of pilot shafts,

and drilling of drawbells). The minimum horizontal distance between the undercutting front and

the development and constructions, in the production level, is frequently the vertical distance

 between the undercut level and the production level. This is called “the 45 degree rule” (Brown,

2007). For example, in the case of the Esmeralda sector of El Teniente Mine — despite the fact

that the distance between levels is 18 meters — the minimum distance used was 22.5 meters

(Jofre, 2000).

2.2.5 Cavability

Brown (2007) defined cavability as “a measure (often non-quantitative) of the ability of an

orebody to cave under particular circumstances” Cavability is associated with a minimum  area

(area of caving) and a minimum span, which is the minimum width of the undercutting front.

This area of caving (hydraulic radius of caving, hydraulic radius is calculated dividing the area

 by the perimeter of undercut zone) has significant importance in the stability of in the

undercutting front, because when the undercutting area achieves this value, the caving process

starts and the roof of the cavity collapses —  which changes the abutment stress around the caving

area.

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Figure 2.6 Illustration of crinkle cut in Palabora mine (Brown, 2007)

Figure 2.7 Crinkle cut undercut potential problem areas (Butcher, 2000)

2.2.6 Undercutting rate

There is not an optimal value for the undercutting rate, as it depends on the characteristics of

the rock mass, the capacity of the development, the construction, and the production

requirements (Flores, 2002). From the geomechanical point of view, the undercutting rate has

influence on the stability of the abutment stress zone (Karzulovic, 2001). If the rate is very low

or the undercutting front stops, the magnitude of the stress begins to open structures and create

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new fractures (Butcher, 1999). This decreases the strength of the rock mass, provoking damage

in the mining infrastructure. On the other hand, if the undercutting rate is too fast, it can originate

high seismicity with the possibility of rock burst (Karzulovic, 2006). An average rate of

undercutting is around 2,250 m2/month (Figure 2.8) (Flores, 2002).

Figure 2.8 Relative frequency of different undercut rates in caving mines (Flores, 2002)

2.2.7 Geometry of undercutting front

The geometry of the undercutting front has an effect on the extent of damage ahead of the

undercut front, basically because it can generate the stress of confinement, in the case of concave

geometry, or traction, in the case of convex geometry. A convex geometry should be avoided

 because it generates more extensive damage (Figure 2.9) (Laubscher, 2000).

Based on Laubscher’s work (1994), Butcher (1999) developed five rules to avoid damage in

 block caving operations. One of the most important is the geometry of the undercutting front.

Butcher points out that the problems associated with the geometry of the undercut are the large

irregularities in the horizontal geometry of the undercut front (cave line), which cause an

increase in stress concentration that may result in serious damage, (Figure 2.10). Therefore, the

horizontal irregularity in the undercut front panel should be avoided (Brown, 2007).

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Figure 2.9 Confinement and traction in the undercutting front (modified from Laubscher,2000)

Figure 2.10 Geometries of the undercutting front (Butcher, 1999)

2.2.8 Undercutting method in high stress

Mining in later years has demonstrated that caving methods can be applied successfully in

strong rock mass conditions at greater depths, which increases geotechnical challenges (Flores,

2008). Together with the deepening of mining operations, there has been an increase in stress

conditions, (Figure 2.11). The use of post-undercutting in a condition of high in-situ stress

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(Flores, 2002) causes significant redistributions of stress (abutment stress) in the caving front,

which causes damage in the production levels that in turn affects the available production area

and increases the number of repairs. Figure 2.12 shows the increase in stress caused by the

undercutting advance effect in the roof of a production drift (Bartlett, 1992).

Figure 2.11 Variation of in-situ stress with depth (Karzulovic, 2004)

Figure 2.12 Stress conditions in the roof of production drift by effect of the undercuttingfront advance (Trueman, 2002)

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Some changes in the sequence of development and construction of mining infrastructure have

 been applied to solve these problems in order to manage the stress. In high-stress conditions,

 postponing all or part of the development and construction in the production levels until the

undercut front has passed can prevent premature damage. This causes a decrease in the width of

the pillars in the undercut level, which leads to a reduction in strength of the pillars, difficulty the

undercutting blasting due to the damage in the blast hole, thus causing partial blasting and

therefore the formation of remnant pillars (Villegas, 2008) (Figure 2.13, Figure 2.14 and Figure

2.15)  or even collapses in undercut and production levels. For these reasons, some mines are

considering returning to the post-undercut method and will assume the costs of reparation, but

will gain flexibility and stability for the undercut level (Araneda, 2007).

Figure 2.13 Example of advanced undercutting front with small pillars (Workshopundercutting, 2008)

2.2.9 Generation of dynamic stress

In all undercutting methods used in block caving, the undercut is performed by drilling and

 blasting to cut the base of the block. This is a discontinue process that increases stress due to the

explosive detonation and the instantaneous redistribution of stress caused by the volume

removed from the base of the block by the undercutting blast (Aguirre, 1994).

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In order to control these effects, every mine carefully plans its undercutting program and

avoids blasting large areas or using significant amounts of explosives (Flores, 2002). Mine

 project staff normally monitors the vibrations and seismic activity to adjust the blasting design.

The effect of dynamic stress increases if the in-situ stress condition of high stress, because it

causes a major transfer of energy around the blasting area (Dunlop, 1999). For this reason, the

 blasting of drawbells can cause minor damage in a pre- or advance undercut method where the

stress condition is relaxed after the passage of the undercut front (Potvin, 2010).

Figure 2.14 Remnant pillar in undercut level (Karzulovic, 2005)

Figure 2.15 Damage caused by collapse in production level (Karzulovic, 2003)

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2.2.10 Seismicity associated to undercutting process

The relationship between the undercutting advance with drill and blast and seismicity has

 been studied in El Teniente mine (Potvin, 2010). The conclusion reached was that there is a

direct cause/effect relationship between the two. To illustrate this concept,  Figure 2.16 presents

the relationship between the undercut area and the seismic events in the sector RENO of El

Teniente Mine (historically known for the rockburst). The red line represents the cumulative

number of significant seismic events (seismic events that the people in the El Teniente mine

could feel) in the timeframe measured. The green line shows the cumulative undercutting area by

drill and blast. Clearly, there is a direct relationship between the frequency of the undercutting

 blasts and the increase in seismicity.

In addition, it is interesting to note that one of the strategies used to control the rockburst in

 block caving is to decrease the area of undercut in each blast along with the frequency betweenthe blasts, to decrease the number and magnitude of the seismic events.

Figure 2.16 Cumulative undercut area blasted (green dots) compared to the cumulativenumber of significant events (red line) showing a very strong correlation(Potvin, 2010)

2.2.11 Numerical models of undercutting in block caving

The most common use of numerical models in block caving is associated to the caving

 propagation relationship with the extraction rate, namely beyond hydraulic radius, and parameter

studies such as:

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-  Geometry of the cutting front during the font advance, in order to study the stability of

undercut and production levels. Also, these analyses allow selecting the suitable,

undercutting method (Brown, 2007).

- The caving rate, in this case the focus is on trying to predict when and how the fracturing

arises in the surface. This is important information in the transition from open pit to

underground method (Brown, 2007).

- The dilution prediction, this is a key economical factor, because it is related to the

mineral recover and it is an input for planning the needs of reposition of the production

area (Board, 2009).

- Subsidence estimation, this prediction is important for the location of infrastructure,

shaft, crushers, and other that can be affected by the caving process. In addition, this

 parameter may be related to with the dilution entry (Board, 2006).

Usually, the numerical models in block caving are three dimensional, because the caving

 process affects the whole rock mass around the caving volume.

There is a significant uncertainty on the transition from intact rock to caving material

(Sainsbury, 2010), the propagation rate of caving and the relation between the rock mass quality

and the stress condition. In general, the numerical model of caving uses strain softening as

constitutive property, which through the evaluation of “critical plastic strain” adjusts the rock

 properties and simulates the rock mass degradation around caving volume.

2.3  Longwall mining method

Longwall mining is a highly mechanized underground cave mining technique that is usually

used to extract a seam from a bedded sedimentary deposit such as coal or trona. Seam heights

are in the range of 2 to 5 m, which is relatively thin compared to block caving column heights

Within non-metallic underground methods, longwall method has the highest productivity

(Haycocks, 1997). The drawback is the high capital cost of equipment and installation (Energy

Information Administration, 1994). The main components in longwall are hydraulic roof support,

a cutting machine, and an armored face conveyor (Figure 2.17) (Mitchell, 2012).

The longwall system advances along a full face wall of rock with a cutting machine, and

loads the coal mineral into a conveyor system parallel to the cutting face. A shield of hydraulic

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support that advances together with the undercutting front advance protects the whole equipment

system. Behind the shield support, the roof collapses, filling the cavity (Mitchell, 2012).

This method is used mainly in soft rock such as coal or potash. A disadvantage of longwall

mining is the high level of skill and training required for personnel operating the system (Energy

Information Administration, 1994). Moreover, if some of the components have problems, it stops

the entire production.

Figure 2.17 Classic schematic of longwall, in a coal mine (Energy InformationAdministration, 1994)

2.3.1 Mine design

A typical layout for a longwall coal mine is given in Figure 2.18. A series of roadways called

main entries service a section of the mine. The roadways are developed using continuous mining

machines (Peng, 2006). These entries are used for the main mine ventilation system, and provideaccess for people, machinery, electrical supply, communication systems, water pump out lines,

compressed air lines and gas drainage lines.

The longwall design is related to the geometry of the coal seam such as height, dip, depth etc.

and the production requirement (Peng, 2006). The geological and geotechnical investigations are

very important to design suitable longwall application and choose the equipment properly.

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Parameters such as water table, aquifers, faults and lithology of the rock columns, may define the

success or failure of the method implementation. During the longwall designs is necessary to

estimate the longwall production and define the width and length of the panel (Energy

Information Administration, 1995), the number of entries, size of pillars, advance rate, advance

direction, ventilation requirements and others.

A panel of coal to be extracted is blocked out using entries along the sides of the panel and

connected together across the end of the panel. The pillars formed between adjacent entries are

called chain pillars. In a developed longwall panel, the headgate is the entry where the belt

conveyor system, the stage loader and crushing system used to transport the extracted coal are

located. The tailgate entries are on the opposite side of the panel and are used to complete the

ventilation circuit to remove methane and dust from the longwall cutting face. A barrier pillar is

often left at the end of the panel to protect the main entries. Panel dimensions range from 150 to300 meters wide, and 1000 to 3500 meters long.

Other typical longwall drift are bleeder entries (allow the ventilation to eliminate the methane

concentration) and the main entries that allow the connection with other panel and access to the

surface (Energy Information Administration, 1994). Longwall could be advancing or retreat type.

In the advancing type the gate roads are formed as the coal face advances (Wikipedia, 2014). In

retreat longwall the panel is formed by driving the headgate, tailgate and development before the

cutting face advances. Figure 2.18 shows a classical longwall retreat design.

2.3.2 Production system

Currently in longwall all the production systems are mechanized. The roadways are

developed using continuous mining machines (Peng, 2006). The production is obtained cutting

the coal seam or other mineral to full face using a cutting machine called  shearer , which consists

in a drum with cutting picks (between 40 to 60) which cuts the rock in a continuous way. The

rock falls over a conveyor system (AFC armored face conveyor ) and is extracted by means of a

 belt system to the surface. The production operation is in the face, to protect the people and

equipment a shield  is used, which consists in hydraulic jacks that support the roof and advances

coordinately; Figure 2.19 shows the production system (University of Wollongong, 2014).

The shearer haulage system is attached to the AFC, through a mechanical interface, the

system pushes the drum to the face and this rotates cutting the coal, also the AFC is connected

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with the hydraulic shield and everything is an interconnected system (University of Wollongong,

2014).

Figure 2.18 A typical longwall coal mine layout showing the panels, entries, and chain pillars, (Source: Ch13.8, SME Handbook, 2011).

Hydraulic shield is the system used to support the roof, this allows the work of the cutting

machine and protects the people and production equipment (Barczak, 1992) (Peng, 2006):

The relationship between the support load, generated by a shield support and the stiffness of

the support in terms of roof convergence, prior to generating the yield load is critical to control

the roof above the work area (Mitchell, 2012). The purpose of hydraulic shield support is to

 prevent or delay the failure of the roof during the mining operation (Barczak, 1992).

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The AFC, is formed for short sections, called pans, which allows to move the material from

the cutting face. The section is 1.5 meters long average, the transport velocity is around 1

meter/second. The AFC is installed in front of the hydraulic support and behind shear system.

All the system advances together.

In longwall, a crusher is used to reduce the coal fragment to the appropriate size for the belt

design and then it is loaded onto the first conveyor belt by the BSL (Beam Stage Loader) (Peng,

2006).

The components of the longwall machine are highly integrated and controlled by a

computerized control system. When the shearer passes, individual shields are advanced forward,

and the entire longhwall system advances thereby together in a systematic fashion.

Figure 2.19 Overview of longwall system, main equipment (University of Wollongongwebsite, 2014)

2.3.3 Numerical model of longwall

The numerical models are widely used in the longwall studies. The main analyses are related

to the stability of gates (Shabanimashcool, 2012), in order to ensure the safety of the workers and

infrastructure stability, also the numerical models are used to evaluate the mine design, panelorientation and effect of undercut advance (Chen, 1998) (Qing- Sheng, 2013). Regarding design

of pillars, chain pillars and yielding pillar, numerical model are used to estimate the strength of

 pillars (Badr, 2002).

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The hydraulic support and shield support, are inherent aspects to the longwall method, in

which is necessary to estimate the load variation during the undercut front advance (Barczak,

1992), namely the stress and deformations, for these analyses numerical model helps to predict

 possible behavior in different rock mass conditions. In order to study the behavior of these

 parameters, the numerical model must incorporate the strength of rock mass beyond the peak

strength, incorporating the residual strength parameters. With these concepts most of the

numerical models have used strain softening as constitutive property (Ozbay, 2003).

On the other hand, the longwall method is a caving method, the mineral is extracted and the

cavity is filled by surrounding material, fracturing advances in vertical direction (Peng, 2006).

Velocity and magnitude of vertical movement depend on the depth of exploitation and the rock

mass properties, causing subsidence as a consequence (Keilich, 2006). In cases when there is

infrastructure in the surface models are used to study the possible effects (Zhang, 2013).Figure 2.20, shows an example of numerical model in longwall to study the abutment stress

around a panel (Wei, 2011).

Figure 2.20 Three dimensional model (Wei, 2011)

The numerical models in longwall method and block caving method during the modeling of

undercut process are similar, only with the difference in the damage caused by the blasting used

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in the undercutting process in block caving, which is normally confined, this changes the rock

 properties estimation around of the undercutting area, Hoek 1998 presents a correction factor

“D” to re present this effect in the rock mass properties (Cai, 2007). Another important difference

is that in the block caving the mineral column is extracted and the draw process must be

modeled.

2.4  Chuquicamata underground project

This work considers using the stress condition and the rock properties of a real case, where

the mechanical undercutting could be applied.

In this case the Chuquicamata underground project information is considered to be

appropriate, because this project is in the construction phase and is possible to evaluate the

option of using a mechanical system to do the undercut.

2.4.1 Chuquicamata mine

Chuquicamata mine is an open pit mine located near Calama, II Region of Antofagasta,

1,650 kilometers north of Chile’s capital, Santiago. It is approximately 2,900 meters above sea

level (masl) (Figure 2.21). Chuquicamata started its operations in 1915, although the high quality

of the ore has been well known for centuries, dating back to the prehispanic cultures that

inhabited the region. The production is 443,381 metric tons of fine copper electrorefined,

electrowon cathodes, and copper concentrates. In 2011 Chuquicamata had 6,527 employees

(CODELCO, 2011). Today Chuquicamata is a big open pit 4.9 kilometers long, 2.9 kilometers

wide and 1.0 kilometers deep Figure 2.22 (Allende, 2009).

Chuquicamata plans to change its mining method from open pit to underground, because the

open pit is approaching to the end of its economic mine life. The Chuquicamata Underground

Project considers recovering around 1,700 million tons of ore with an average grade of 0.71%

Cu, 499 ppm of Mo and 460 ppm of As, over an operational period of approximately 45 years.

The studies to date suggest that the earliest start of production for the underground mine would

 be towards the end of 2018 at a rate of 4,000 tpd (metric tons per day) of ore from tunnels

development. The production will increase in 2019 by supplementing with ore from the

 beginning of caving activities, followed by a production ramp-up to achieve the design capacity

of 140 ktpd (Fuentes, 2009).

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Figure 2.21 Chuquicamata Underground Mine, location (Fuentes, 2009)

Figure 2.22 Chuquicamata mine (Fuentes, 2009)

The project has considered four exploitation levels, starting in the central area and then

advancing north and south. Two blocks were planned to test new technologies. These blocks will

 begin together with the regular exploitation blocks. This thesis uses the information from one of

those blocks, block N 42 (Figure 2.23), for the different studies.

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Figure 2.23 Macro blocks and geology, first level, Chuquicamata Underground Project(Villegas, 2008)

2.4.2 Geology

The mineralized area of the Chuquicamata deposit is found in a tertiary intrusive body in theimmediate footwall of the West Fault. This intrusive unit corresponds to a porphyric stockwork

of granodiorite composition, known as the Chuquicamata Porphyry. This unit has various

textural and hydrothermal alteration differences that allow it to be subdivided into various units,

including sericitic, potassic, and chloritic domains (Figure 2.24). The mineable portions of the

ore body are found in the sericite-altered rock mass within the immediate footwall of the West

Fault. The highest ore grades (>1% Cu) are found directly adjacent to the fault in the highly

altered and weak RQS, QS) and porphyry (Porphyry East Sericitic, Porphyry East Potassic (K),

and Porphyryc East Choritic) units. The immediate hanging wall of the West Fault is a heavily

sheared, poor-quality Fortuna Granodiorite that increases in rock quality with distance away

from the West Fault (Board, 2006).

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Figure 2.24 Ore deposit characterization (Fuentes, 2009)

2.4.3 Mining method

The mine exploitation method being used is the block caving method, in four levels (Figure

2.25). Two levels will be simultaneously in operation during the production (Fuentes, 2009).

When the underground mine starts, the open pit operation should be stopped. There is not

consideration of a transition time with both operations working.

In this project, the block caving method has been modified to the variant called “macro

 block” (Figure 2.26). This mining method offers advantages in cost and production capacity

(Fuentes, 2009). The characteristic of this method is to separate and group the activities of

developments and construction in one sector without interferences from other activities.

The advantages of macro block method with respect to panel caving are:

  It gives more flexibility in production planning, developments of mining infrastructure,

and other operations.

  It is suitable for the geometry of the Chuquicamata’s ore body (long and thin).

 

The modular design allows the incorporation of technological changes more easily.

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  Major collapses in the production areas could be isolated easily.

 

The activities of development, construction, and undercutting are separated from the

 production operation with the possibility of achieving a better index of productivity.

Figure 2.25 Position of production levels in Chuquicamata Underground Project(Fuentes, 2009)

Figure 2.26 Macro block method (Fuentes, 2009)

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The macro block method was developed during the Chuquicamata Underground Project as a

way of managing mining preparation and production administration. It is a variant of block

caving, but with larger block size and with a pillar to separate the macro blocks. Its configuration

was developed by the Chuquicamata Underground Project to confront the important need of

 preparing and replacing the depletion area in order to keep production over 140,000 tons/day

(Fuentes, 2009). With this configuration, it is possible to isolate zones that have stability

 problems such as collapses, air blast risk, and mud rushes without stopping new areas from being

incorporated into the production. The macro block method is a variant of block caving that offers

great management opportunities to organize the activities of construction and production. Each

macro block could be considered an independent unit with its own mine design and organization.

Figure 2.27 shows the macro block configuration for the Chuquicamata Underground Project

(Villegas, 2008). The method gives flexibility to incorporate new technologies or adjust designs.Each macro block should have enough area to initiate a new caving. In addition, each new macro

 block should have a special draw system, to control the dilution from neighboring blocks

(Fuentes, 2009).

Figure 2.27 Macro block configuration in first and second levels of Chuquicamata UndergroundProject (Allende, 2009)

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CHAPTER 3

UNDERCUTTING WITH A MECHANICAL SYSTEM

3.1  Block caving mine designs with mechanical undercutting

This work presents an alternative to the use of the drill and blasting method for the

undercutting process; in this case, using a mechanical system similar to the one used in a

longwall panel in underground coal mining. For this idea to be feasible, the undercutting process

should be compatible with the production process. Therefore, it is necessary to have a mine

design that incorporates both activities.

In the longwall panel method, the undercut front advances as the coal is cut and removed

and behind it, the overlying rockmass collapses and is left behind, this area is called the Gob. In

the block caving case, the Gob area is above the production level, where the caved rock isremoved through drawpoints.

One important aspect to consider for the implementation of this mechanical undercutting

system in block caving is the undercutting process, because when undercutting in block caving is

developed, in any of its forms, the undercutting activity is located ahead of the cutting front, but

in a mechanical system such as longwall is located behind the cutting front — namely, inside of

the caving, Figure 3.1. 

Figure 3.1 Differences in location of the work area in classical undercutting andmechanical undercutting system

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When applying a longwall like mechanical undercutting system in block caving, the

cutting process is accomplished from the inside of the undercut area. Therefore, when using a

 block caving exploitation system, it is necessary to separate the undercutting zone from the

 production zone or caving volume (where caving grows by the extraction process) Figure 3.2. 

This is possible using pre or advance undercut. In this configuration, the cutting system advances

along with the undercut, and the drawbells are opened in the back to incorporate the area into the

caving. For this reason, it is important to understand what might be the effect of the caving

advance in the undercutting process, which is studied in Chapter 5.

Figure 3.2 Scheme of mechanical undercutting system in block caving (modified fromBrown, 2007)

3.1.1 Undercutting mine design

The layout of the mine design for a mechanical undercutting system should consider the

following requirements.

Access to the cutting front is needed for:

o  The assembly and disassembly of the undercutting machine,

o  Access to the work face for operation and maintenance of equipment,

o  Ventilation of the front face, and

o  Handling of muck.

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Separation of undercut panels is needed for:

o  Allowing access to the next panel to be undercut

o  Use of all the potential associated with the block caving method by macro blocks.

Dividing a panel to reduce the width of the undercutting front provides:  

o  Control of the stability in the work area, and

o  The ability to increase the rate of the undercutting front advance.

The design proposed as shown